Flotation with xanthate

ABSTRACT

A process for the flotation of copper minerals employing xanthates, wherein the copper minerals are strongly activated in at least one stage at a pH of 10 to 12 and then conditioned with a dispersing agent or a flocculating agent prior to flotation, the process showing marked improvement in grade and recovery in the case of refractory ores and ore containing talc and clay, and being effective to concentrate copper sulphides, copper silicates and copper oxide in the same circuit.

United States Patent 1191 Weston 1 1 FLOTATION WITH XANTHATE David Weston, 34 Parkwood Ave., Toronto, Ontario, Canada [76] Inventor:

[ Notice: The portion of the term of this patent subsequent to May 29, 1990, has been disclaimed.

22 Filed: Nov. 6, 1972 211 App]. No.: 303,691

Related U.S. Application Data [63] Continuation of Ser. No. 874,026, Nov. 3, 1969,

abandoned.

[52] US. Cl. 241/24; 209/5; 209/166 [51] Int. Cl. B02c 17/00 [58] Field of Search 209/l66,l67, 5; 241/20, 241/24 [56] References Cited UNITED STATES PATENTS 1,286,532 12/1918 Christensen 209/166 X 1,417,263 5/1922 Luckenbach 209/166 1,659,396 2/1928 Douglass 209/166 1,973,278 /1934 Barker 209/166 1,973,578 9/1934 Ruth 209/166 2,310,240 2/1943 Keck 209/166 2,349,637 5/ 1 944 Ruckwardt. 09/ 1 67; l 66 2,469,368 5/1949 Carvajal 209/166 2,471,384 3/1944 Booth 209/166 2,485,083 10/1949 Booth 209/166 2,628,716 2/1953 H111 209/166 2,628,717 2/1953 Booth 209/166 2,901,167 8/1959 Fischer 209/166 3,220,551 11/1965 Moyer 209/167 3,252,662 5/1966 Lyons 209/166 X 3,375,924 4/1968 Corbett... 20 /167 3,386,572 6/1968 Cadwell 209/166 X 3,456,792 7/1969 Schoolcraft 209/167 3,528,784 9/1970 Green 209/167 X 3,539,002 11/1970 Last 209/167 X 3,596,838 8/1971 Weston... 209/166 X 3,667,690 6/1972 Weston... 209/166 X 3,735,931 5/1973 Weston 209/166 X OTHER PUBLICATIONS Chem. Abst., 65, 1966, 10194h, 10195a.

Chem. Abst., Vol. 167, pp. 92957b, 929581. I Froth Flotation, 50th Anniversary Vol., 385, 390, 391, 408, 409 1962.

Froth Flotation, 50th Anniversary Vo1., 382384, 386-389, 392-400.

Primary ExaminerRobert Halper Attorney, Agent, or Firm-Depaoli & OBrien [5 7] ABSTRACT 15 Claims, No Drawings FLOTATION WITH XANTHATE This is a continuation, of application Ser. No. 874,026 filed Nov. 3, I969 now abandoned.

BACKGROUND OF THE INVENTION This invention relates to improvements in the flotation of copper minerals wherein xanthate and the related families of chemicals that are known to act as sulphide collectors (hereinafter referred to collectively as xanthates) are used as collecting agents.

Certain copper ores have not heretofore responded satisfactorily to flotation with xanthates. Some of such ores have conventionally been treated with sulphidizing agents such as sodium polysulphide and sodium sulphite or have been treated with combinations of xanthates and other collectors such as fatty acids. Even with the employment of such measures the grade and recovery achieved have not been satisfactory but have had to be accepted in the treatment of these so called refractory ores. The copper minerals in such ores frequently are associated with substantial portions of clay or talc which give rise to additional difficulties in their treatment by conventional flotation techniques.

It is the object of the present invention to provide a process for the flotation of such ores with xanthate collectors giving marked improvement both in recovery and grade compared to previously employed flotation processes.

It is a further object of the invention to float, in the same circuit, with high recoveries, such widely diverse copper sulphides as Chalcopyrite, Bornite and Chalcocite and the so called acid soluble copper minerals such as Chryscolla, (copper silicate), and Malachite and Azurite, (the common carbonate copper minerals).

It is a further object of the present invention to carry out such flotation to produce high grade of concentrate and high copper mineral recoveries regardless of the presence in the ores of such deleterious materials as talc and clay.

SUMMARY OF THE INVENTION With these and other objects in view the invention broadly stated consists in conditioning, in at least one stage, an appropriately prepared pulp of an ore containing copper minerals in the presence of an alkaline agent and a xanthate at a pH of from about to about 12 to produce heavy activation of the copper minerals; then subjecting the pulp to a final stage of conditioning in the presence of a dispersing agent or a flocculating agent until the host rock ingredients of the ore are depressed; and then subjecting the pulp to flotation to produce a concentrate of the copper minerals.

The expression appropriately prepared pulp of an ore" when used herein is intended to mean that the pulp has been made up from ore which has been comminuted to the extent which by conventional standards would be considered sufficient for acceptable liberation of the desired mineral constituents and has during such comminution or thereafter been subjected to such treatment steps (such as adjustment of pulp density, and adjustment of pH) as the operator may deem appropriate, in the case of the particular ore being treated, to present the .pulp for the stages of treatment comprising the process of the present invention. These treatment steps may be conventional and for conventional purposes, or they may, in conjunction with the basic process of the present invention, form one or more stages of a preferred embodiment of the invention.

In the preferred embodiments of the invention the comminution circuit may be considered the first stage as the inventor has found thatfor optimum results initial pH control in the comminution circuit is important, as is also the reagents used for pH control, even though the pH in this stage may be radically different from the optimum pH of the flotability of the minerals.

Further, in variations of the invention it may or may not be beneficial to add the collector to this circuit.

In other variations, the first conditioning cycle may simply be a continuation of the controlled conditions established in the pulp as the final product leaves the comminution circuit.

By alkaline agent is meant any agent used for upward adjustment of the pH of the pulp and which is non-deleterious to flotation.

By the expression heavy activation" is meant a state of activation in which at least the desired minerals are so heavily activated that they will float in a much shorter time than in current conventional practice, and in addition will not be depressed in the cleaner circuit wherein the pH may be appreciably higher than the pH used in the rougher circuit, and appreciably higher than the accepted optimum pH for flotation of the desired recoverable minerals. Heavy activation does not preclude substantial activation of minerals other than the copper minerals and the expression is used to distinguish from the present day conventional practice wherein the activation of the minerals is accomplished by stage addition during flotation of much smaller amounts of xanthate with a view, over the complete flotation cycle, to activating more and more of the desired mineral in stages while ideally activating none of the other minerals constituting the ore.

By dispersing agent is intended to be meant an agent which is non-deleterious to flotation and which tends to create conditions of dispersion within the pulp at the prevailing pH.

By flocculating agent is intended to be meant an agent non-deleterious to flotation which tends to create within a pulp, at the prevailing pH, conditions of flocculation.

Preferably according to the invention, following the rougher float, the pulp is subjected to a further period of conditioning in the presence of a dispersing agent or flocculating agent to further depress the host rock ingredients of the ore and the pulp is then subjected to a first cleaner float. Additional cleaner stages may be desirable in many instances and various optional procedures may be followed in the subsequent cleaning stages. On the particular ores treated in the following examples it has been found that best results are obtained by employing the above mentioned conditioning stage between the rougher float and the first cleaner float.

The copper ores upon which the process of the present invention has proved most successful contain acid soluble copper minerals and in the inventors circuits various amounts of copper go into solution. I have found, surprisingly, that the tendency of the copper to go into solution may be minimized and the recover of copper in the concentrate correspondingly increased if the pulp is prepared in a particular way. Thus, appreciable lowering of copper in solution occurs if the ore is ground in an acid circuit for instance at a pH below about 5.5 prior to the raising of the pH to the range of -12 for activation, and furthermore the effect is enhanced if, following grinding, the pulp is given a stage of conditioning with a xanthate at an intermediate pH of from about 6 to about 9.5.

DESCRIPTION OF THE PREFERRED EMBODIMENTS A more detailed understanding of the invention may be had by reference to the following examples of laboratory testing on three different types of ores occurring in the same major copper deposit in the United States of America. These ores are identified as follows:

Lot 130 This sample was a composite of daily concentrator feed samples, and could be described as medium in its clay and talc composition.

Cu (Total) 1.09% Cu (Acid-Soluble) 0.36% Fe 7.55% S 3.23%

The ore body from which these samples were taken is that of a major copper producer in the United States and the producers standard laboratory tests gave the following metallurgy. All three samples contain varying amounts of talc and/or clay.

Lot 128 Concentrate "/1 Cu 13.2%

% of Total Cu in Concentrate 75.8%

Rougher Tailing 7: Cu 0.24%

Lot 126 Concentrate "/1 Cu 8.85%

'7'- of Total Cu in Concentrate 74.9371

Rougher tailing '71 Cu 0.35%

Lot 130 Concentrate 7( Cu L537! 7: of Total (u in Concentrate 76.567: 0.27%

Rougher Tailing 71 (u The foregoing may be considered as indicative of the best metallurgy obtainable on these ores in accordance with the established practice of this operator and the recoveries indicated were accepted as the maximum possible recovery for these ores at the concentrate grades obtained.

A surprising thing concerning the process of the present invention is that it produces both an increase in grade and an increase in recovery.

In the examples, in all cases but one grinding was carried out on 800 gram charges in a laboratory rod-mill to a screen analysis of approximately mesh, and at a pulp density of 50% solids. The pulp was transferred directly from the rod mill to a 600 gram laboratory Fagergren cell, and all conditioning and flotation was carried out in this cell. No desliming stage was used.

All cleaning steps were carried out in a Denver 250 gram cell.

Except where noted in the specific tests the rougher flotation time was seven minutes, and in the cleaners the float was continued in each case to final clean-up with no set time. The average cleaner float time was about four minutes. The rougher float time is exceptionally low, as with this class of mineral, for any degree of recovery in a conventional circuit, a minimum float time of 15 minutes would be required.

With an 800 gram charge to a 600 gram Fagergren cell, the density of the pulp was approximately 33% solids. This is an unusual feature of the invention in view of the tale and clay contents of the ores treated. On normal ores, without talc or clay minerals present, the flotation density in current plant practice is approximately 25% solids. At this low density, for the same flotation time, more cells would be required not only raising capital plant costs but also operating costs in power and maintenance.

In the following series of tests the rougher tailing and first cleaner tailing are shown only for comparative metallurgical reasons, and where the ratio of concentration at this point in the circuit is in the range of about 10 12 to 1. In all cases, on the three samples, complete circuits are described illustrating the production of concentrate grades in excess of 20% copper.

The copper analysis was done by the current standard atomic absorption method with a number of the testing programme samples check analysed by the supplier of the various ore samples. Where the term corrected is used, this indicates application of the correction factor necessary to relate the laboratory analysis to the check analyses of the supplier. To the combined rougher and first cleaner tailing, to correct to the suppliers analysis, the rule of thumb correction factor is plus 0.01% copper.

In both an all alkaline circuit and a circuit where a preliminary acid stage was used an average of 2 /2% of the total copper reports in solution. Except in the tests showing the solution analysis, all analyses are on the solids only.

Lot 128 High Clay Content EXAMPLE I (TEST 8) The sample was ground with 8.75 lbs. per ton of sodium carbonate producing a pH of 9.85. The pulp was then conditioned with 0.15 lbs. per ton of Potassium Amyl Xanthate (Z6) for 15 minutes.

20 lbs. per ton of sodium silicate was then added to the pulp for 5 minutes together with 5 drops of pine oil for the last 3 minutes. The final pH was 10.0.

The rougher float was conducted for 7 minutes with the addition of 0.025 lbs. per ton of Z6 and one drop of pine oil.

The rougher concentrate was conditioned in the Denver Cell for 3 minutes with 0.025 lbs. per torrof'Z and 2 lbs. per ton of sodium silicate and floated to clean-up.

The rougher plus cleaner tailings was 91.8% by weight analyzing 0.77% copper. In comparison to the conventional test where the rougher tailings loss was 0.24% Cu, this represents a major increase in recovery.

A surprising feature is that the optimum pH for flotation of this ore was previously considered to be 11.6 whereas this float was made at a pH of 10.0.

EXAMPLE 11 (TEST 5) This test was identical to Example 1 with the exception of lime replacing the sodium carbonate in the first stage to bring the original pulp during conditioning with the collector to the considered optimum pH. The pH at the end of the conditioning with Z6 (potassium amyl xanthate, American Cyanamid) was 11.65. However, following the addition of the strong alkali dispersant, sodium silicate, the pH immediately dropped to l 1.1 and the rougher flotation was carried out at a pH of 10.9. Why this drop in pH occurs is not known.

The rougher plus first cleaner tailing was 91.8% by weight analyzing 0.72% Cu. One of the surprising results of this test was that even though the activation cycle with Z6 was at the considered optimum pH and the final pH was still much closer to the optimum than in Example 1, the metallurgy was nearly identical.

This indicated that the previously optimum flotation pH for these minerals was no longer a critical factor in the new circuit. In other words the preliminary stages of the process are more important than the precise final pH in the flotation circuit.

Although the lime is a flocculant, in comparison to Example 1 where soda ash, a dispersant was used, there was no significant difference in final metallurgy.

EXAMPLE lll (TEST 1) To the grinding stage 5.0 lbs. per tone of lime and 5.0 lbs. per ton of soda ash were added resulting a pH of 11.4. 0.20 lbs. per ton of Z6 was added and the pulp was conditioned for 30 minutes. Following this stage 0.75 lbs. per ton of a flocculant was added (a locust bean gum). The pulp was conditioned for seven minutes with the final addition of pine oil as a frother. The rougher float was eight minutes with the addition of 0.025 lbs. per ton of Z6 and one drop of pine oil. The cleaner used four minutes of preconditioning with 2.4 lbs. per ton of soda ash, 0.025 lbs. per ton of Z6, and 0.25 lbs. per ton of locust bean gum. The combined rougher and first cleaner tailing was 90.6% by weight analyzing 0.09% Cu. This was surprisingly good metallurgy for the substitution of a flocculant for this cycle in place of a, dispersant, and although the metallurgy was poorer than Examples 1 and 11, it was appreciably better than the conventional circuit.

This test showed that where an all lime circuit was not used in the initial stages, in the final stage either a dispersant or flocculant could be applied.

EXAMPLE IV (TEST 34) r! In this example 0.5 lbs. per ton of soda ash was added to the grinding stage with the pH at the end of the grind being 6.45. To the first conditioning stage 6.8 lbs. of lime and 0.1 lbs. of NaCN per ton respectively were added for a conditioning time of 12 minutes. The pH at the beginning of the cycle was 1 1.8. The cyanide was added for pyrite depression. The next cycle was minutes conditioning with 0.17 lbs. of Z6 stage added.

1n the last stage before floating 7.5 of sodium silicate per ton was added and conditioned for 5 minutes with three drops of pine oil for the last three minutes.

The rougher float was 7 minutes with the addition of 0.025lbs. of Z6 per ton and one drop of pine oil. The first cleaner float used 4 minutes pre-conditioning with 0.375 lbs. oflime, 0.05 lbs. of sodium cyanide, and 1.25 lbs. of sodium silicate per ton respectively.

The second cleaner float used 6 minutes preconditioning with 1.0 lbs. lime, 0.10 lbs. sodium cyanide and 0.5 lbs. sodium silicate per ton respectively.

the following corrected metallurgy was obtained.

Product Wt. Analysis Cu Distribution Concentrate 3.3 21.6 79.8 No. 2 clnr tlg. 1.1 2.07 2.5 No. 1 clnr tlg. 7.0 0.48 3.8 Rougher tlg. 88.6 0.14 13.9

than 50% higher with an increase in copper recovery,

even though the use of the cyanide as a pyrite depressant had not been optimized.

LOT 126 High Talc. High Pyrite Content EXAMPLE V (TEST 37) 1n the grinding stage 2.5 lbs. per ton of soda ash was used with the pH at 7.0.

7.0 lbs. oflime, and 0.10 lbs of sodium cyanide were added to the next stage for 12 minutes conditioning with a pH of 12.0 at the start and 11.85 at the end of the cycle.

In the next cycle 0.17 lbs. of Z6 per ton was stage added and conditioned for twenty minutes. The end pH was 1 1.75.

1n the final cycle before flotation 7 /2 lbs. per ton of sodium silicate was added and 3 drops of pine oil for the last 3 minutes. The conditioning period was 5 minutes. The rougher flotation was 12 minutes with the addition of 0.025 lbs. Z6 per ton and one drop of pine oil.

The first cleaner float used 6 minutes preconditioning with 0.75 lbs. of lime, 0.075 lbs. of sodium cyanide and 1.25 lbs. of sodium silicate per ton respectively.

The second cleaner float used eight minutes preconditioning with l.25 lbs. lime, 0.10 lbs. of sodium eyanide, and 0.625 lbs. of sodium silicate per ton respectively.

, The following metallurgy was obtained:

sultcd in only low drops in copper values in the cleaner tailings. This test illustrates the heavy activation" of the copper minerals that was obtained prior to the rougher float.

crease in copper recovery.

EXAMPLE Vl (TEST 42) Lot 130 The following metallurgy was obtained using cya- EXAMPLE nide, and lime in the cleaners. The chemical analyses l of the various products are corrected and were as fol- A Sales of tests was dlfferejt alkdlme lows; agents to control the pH with all other variables kept constant while adjusting the pH to the optimum" of Aflnfllysis 1 1.6 in the activation cycle with xanthate. Product Wt. Cu Distribution The circuit was as follows:

Concentrate 2.5 35.7 83.5 No. 3 clnr tlg 2.7 2.1 5.3 NO. 2 r g Grinding Stage twelve minutes with alkali No. l elnr tlg 12.6 0.19 2.3 agents Show in COL 1 Rougher 0067 First Cycle two minutes conditioning with lime 100.0 100.0 bringing pH to approximately l 1.6

- Second Cycle nineteen minutes conditioning with stage addition of Z6 Third Cycle five minutes conditioning with sodium Not only is the grade of concentrate increased by in silicate and pine oil excess of 100% over the standard but also the open cir- "9 Stage addmm (W cuit recovery in copper values is appreciably increased, one drop of pine oil lst Cleaner three minutes conditioning with 0.025 even before taking into consideration the add tional 25 lbs 26 and L25 lbs Sodium Silicate copper recovery that would be made on closed clrcuitrespectively f ll d by ing of the second and third cleaner tails. "Malian Clean-U11 pH Rghr. Plus Clnr. Tlgs. Test Reagent to Reagent After Z6 "/0 Wt. Analysis 70 Cu No. Grinding Stage Lbs/Ton Grind Cycle 8:. Test No. 1 Lime 40 8.75 11.6 11.6 92.0 0.099 2 Lime 44 0.5 7.9 11.6 91.3 0.072 3 Soda Ash 42 0.625 7.55 1 1 6 92.0 0.073

Potassium 4 Carbonate (43) 0.625 8.4 l 1.6 91.4 0.091

Sodium 5 Hydroxide 45 0.5 8.0 1 1.55 1.1 0.081

Ammonium n Hydroxide (46) 0.5 7.5 11.52 91.4 0.076

The following circuit was used. To the grinding stage 0.625 lbs. per ton of soda ash was used to produce a pH of 7.55. The first conditioning cycle was 2 minutes with 8.5 lbs. of lime per ton resulting in a pH of l 1.75. The second conditioning cycle was 19 minutes with 0.225 lbs. Z6 per ton added in stages. The pH at the end was 1 1.6. The third and last conditioning cycle before flotation was 5 minutes with 6.25 lbs. per ton of sodium silicate and three drops of pine oil for the last two minutes. The sodium silicate dropped the pH to 1 1.4

The rougher float was seven minutes with the addition of 0.025 lbs. of Z6 per ton and one drop of pine oil.

Following the first cleaner, the cleaner concentrate was filtered and re-ground with lime to produce a pH of 12.15.

Following the re-grind stage the pulp was placed in the 250 gram Denver cell and conditioned with 0.1 lbs per ton of NaCN and 0.625 lbs. per ton of Na SiO In the third cleaner the pH was raised to 12.35 with lime and 0.10 lbs. of NaCN per ton of ore and the pulp was conditioned before floating.

On both the second and third cleaners it was surprising that using a pH much higher than the optimum re- It will be observed from the foregoing results that the pH at the end of the grinding is significant and that the copper recovery was significantly better when the pH at the end of grinding was considerably lower than the optimum pH for flotation. Normally one would expect the opposite to be true.

EXAMPLE v111 A series of tests was run which included an acid conditioning cycle prior to the activation of the copper minerals with the collector.

In each case, after the acid cycle, the pH was increased with lime to give a pH of approximately 1 1.6 I

in the activation cycle with xanthate.

The liine cycle was 2 minutes, followed by 19 minutes conditioning with 0.225 lbs. Z6 stage added, and this cycle was followed with the addition of 7.5 lbs. per ton of sodium silicate for 5 minutes with three drops of pine oil added in the last 2 minutes. The rougher float was 7 minutes with the addition of 0.025 lbs Z6 per ton and one drop of pine oil.

The first cleaner was identical in all cases as to reagent balance and time of conditioning.

The following table shows the comparative results.

Grinding Circuit First Cycle Rghl. lst clnr Tlgs Test Reagents Added pH Reagents Added pH Analysis No. & Lbs/Ton End & Lhsflon Begin End 71 Wt. Cu

Soda Ash (47) 7.7 H 80 4.7 5.4" 90.l 0.057 I. 0.625 Lbs. 2.5 lbs Lime (48) 8.0 H SO 4.6 5.5" 90.8 0.065 2. 0.20 lbs. 2.5 lbs H 50 (49) 5.5 Nihil 5.5 5.5 90.9 0.050 3. 2.5 lbs..

It will be noted that with the small amount of sul- EXAMPLE X1 phuric acid added to the grinding stage there is a pronounced reduction in tailings loss. It was at first thought that this could be due to more acid soluble copper going into solution. However, later work did not substantiate the theory. It was a marked improvement in the floated copper minerals.

It will be noted in Test 3 that there was no change in 1 LOT 130 EXAMPLE 1x The third test in Example VIII using an acid stage in the grinding circuit with ten minutes conditioning following the grinding stage at pH of 5.5 resulted in a rougher plus first cleaner tailing rejection of 90.9% by weight analyzing 0.050% Cu. In this Example, following the acid cycle, the pH was raised to the optimum, approximately 1 L6, with lime and the dispersing agent was added at the same time followed by the xanthate activation cycle and then flotation. The tailing rejection was 91.25% by weight analyzing 0.062% Cu. or a small increase in tailings loss compared to the circuit which added the dispersant following the xanthate activation cycle.

EXAMPLE X In this example the sodium silicate dispersing stage was introduced following the grinding stage. To the grinding stage 0.5 lbs per ton of lime was added resulting in an end pH of 7.9. The pulp was then transferred to the Fagergren cell and conditioned for one minute with 6.25 lbs. per ton of sodium silicate. The addition of the sodium silicate raised the pH of the pulp to 8.9. In the next cycle 8.25 lbs. per ton of lime was added raising the pH to l 1.9. In the next cycle 0.225 lbs. per ton of Z6 was added and the pulp conditioned for 19 minutes with three drops of pine oil added for the last three minutes. The end pH of the pulp was 1 1.55. The flotation cycle was seven minutes with the addition of 0.025 lbs. per ton of Z6 and one drop of pine oil. The first cleaner was the standard used in previous tests that is three minutes conditioning with 0.025 lbs per ton of Z6 and 1.25 lbs. per ton of sodium silicate. The tailings rejection was 92.25% by weight analyzing 0.089% Cu. In comparing this test to Example II in series 7 wherein the only change was in the position of addition of the sodium silicate or dispersing cycle it will be noted that the tailings in this test was 91.3% by weight analyzing 0.072% Cu. In other words, addition of the sodium silicate prior to activation with the collector resulted in higher tailings loss.

This example is a duplicate of the third test in series 8 with the exception that following the conditioning stage at a pH of 5.5 the pH was raised with lime to 7.7 and the pulp then dispersed with sodium silicate prior to the raising of the pH of 1 1.6, followed by activation of the copper minerals with xanthane. The tailings rejection in this test was 91.3% by weight analyzing 0.065% Cu as against the tailings loss in the comparative test of 0.050% Cu again illustrating the effect of v the use of a dispersing stage prior to the activation cycle with xanthate.

EXAMPLE XII stage.

EXAMPLE XIII (Test 58) In this example the xanthate collector was added to the grinding stage but the test is otherwise a duplicate of test 2 of Example VII. The tailings rejection was 91.1% by weight analyzing 0.062% Cu showing an improvement in recovery over the addition of the collector in the second last cycle of the comparative test. In this example the contact time of the xanthate collector with the ore was a total of 36 minutes and this gave the lowest-tailings loss of the various all alkaline circuits used.

Contrariwise, where the grinding stage was carried out in an acid circuit the addition of the collector to the grinding stage resulted in lower Cu recovery than in the case where the xanthate was added following conditioning in the acid stage and raising of the pH to the alkaline side.

EXAMPLE XIV (Test 60) This test is a duplicate of the second test in Example VII with the exception that the sodium silicate was replaced by tetra sodium pyrophoshate. In the test using sodium silicate 6.25 lbs per ton of sodium silicate was used as the dispersant. In the case of the pyrophosphate 0.625 lbs per ton was used or only 10% of the sodium silicate. The tailings rejection was 91 .9% by weight analyzing 0.073% Cu. It will be seen that the metallurgy is practically identical showing that the family of phosphates can be effectively used in place of silicates, it being a matter of economics which would be applied.

The silicates would seem to have one major advantage over the use of thc phosphates in plant practice. The surprising factor in the use of sodium silicate was that in the pH range of 10 to 12, as soon as the effective amount was reached there is a sudden drop in pH which can be put to use by automatic controls while with the pyrophosphate there was no change in the pH.

EXAMPLE XV (Test 61) This is a duplicate of Example XlV wherein sodium lignin sulphonate was substituted for sodium silicate or pyrophosphate in the dispersal cycle. The tailings rejection was 90.9% by weight analyzing 0.075% Cu which tailing, since the lignin sulphonate was not optimized, can be considered equivalent to both the sodium silicate and pyrophosphate as dispersing agents. In this example 0.625 lbs per ton of lignin sulphonate was used. The change in pH using the lignin sulphonate was a drop of 0.05. However, as the natural pH of the 5% lignin sulphonate solution used was 8.1 this lower pH can account for the change.

EXAMPLE XVI (Test 64) During the grinding stage of this example 3 /2 lbs per ton of sulfuric acid was added resulting in an end pH of approximately 5.1. Following the grinding stage the pulp was conditioned for ten minutes with no further reagent addition. There was no change of pH during this cycle. At the end of minutes 2.0 lbs of lime and 0.225 lbs of Z6 per ton respectively were added to the pulp raising the pH to approximately 6.5. The pulp was conditioned for ten minutes and at the end of this period lime was added bringing the pH of the pulp up to 11.55 and the pulp was conditioned in this cycle for nineteen minutes. The end pH was 11.45. Following this cycle 12.25 lbs per ton of sodium silicate was added together with three drops of pine oil for the last two minutes of the 5 minutes conditioning cycle. When the sodium silicate addition was made the pH of the pulp immediately dropped from 11.45 to 11.15. The flotation stage was the same as that used in the previous examples. The tailings rejection was 89.4% by weight analyzing 0.048% total Cu and 0.013% acid soluble Cu. Samples of the total solution from the rougher and first cleaner tailings was analyzed for copper and contained 2.2% of the total copper (6.7 of the acid soluble copper). As the solid tailings contained 3.2% of the acid soluble copper the recovery of the acid soluble copper in the cleaner concentrate was 90.1%.

EXAMPLE XVll (Test 65) This example is a duplicate of the previous example with the exception that following the first acid conditioning stage the pH of the pulp was raised to 8.0 and it was conditioned with the xanthate at this point for 10 minutes. The tailings rejection was 90.5% by weight analyzing 0.051% total copper and 0.012% acid soluble copper. The acid soluble copper in solution is equal to 7.6% of the total acid soluble copper, and in the solids tailings is 3.0% for a total of 10.6% so that 89.4% of the acid soluble copper is in the cleaner concentrate prod- UCt.

EXAMPLE XVlll This was a duplicate test of Example Xlll with the exception that following the grinding stage a ten minute conditioning cycle was inserted at a pH of approximately 7.7. The tailings rejection was 90.8% by weight analyzing 0.061% Cu which is practically duplicate metallurgy to the prior example without the use of the additional conditioning cycle. The soluble copper con tent of the rougher plus first cleaner tailings solutions was 10% of the total acid soluble Cu. It will be noted that the acid soluble Cu in solution in this alkaline circuit was appreciably higher than in the prior two examples wherein the original stages were in an acid circuit. This phenomenon is at present unexplained.

EXAMPLE XlX (Test 63) For this example Lot 128 ore was used and a 1000 gram sample was ground for 12 minutes at 50% solids with two lbs per ton of sulphuric acid. After ten minutes of conditioning following grinding without the addition of reagents the pH was 4.5. The pH was raised to 1 1.6 with the addition of 5.65 lbs per ton of lime and 0.225 lbs of Z-6 were added for a 19 minute conditioning period at the end of which the pH was 1 1.45. 12.25 lbs per ton of sodium silicate was added together with three drops of pine oil and the pH dropped to 1 1.10. A flotation cycle similar to that of Example XVI, but carried out at 41% solids, produced a combined rougher plus first cleaner tailings containing 88.4% by weight of the solids and analyzing only 0.042% of the total copper, which is most remarkable metallurgy for such a high pulp density.

As previously mentioned in the foregoing tests about 2 /2% of the total copper reported in the tailings solution. In order to filter the tailings it is helpful to flocculate them by making them acid with sulphuric acid. Surprisingly although the total time involved from the acidulation of the flotation tailings in the flotation cell to the completion of filtration was 2 minutes or less I have found that this copper actually goes into solution during that period. As the following tests will illustrate the tailings solution at the end of flotation in the alkaline circuit contains no dissolved copper at all.

By employing the tailings filtrate as make up solution.

for the grinding of fresh feed material in the acid circuit it is possible to recover most of this sulphuric acid soluble copper by flotation since when the product of the wet grinding circuit is made alkaline the dissolved copper completely reprecipitates and floats with the copper sulphides.

This phenomenon presents a simple and effective manner for rendering sulphuric acid soluble copper minerals amenable to xanthate flotation regardless of whether or not the ore contains clay or talc slimes rendering necessary the use of dispersants and flocculatns in the manner already described.

It appears that for effective flotation of the combined acid soluble and sulphide copper minerals present in ores such as those to which reference has been made it is not necessary that all of the acid soluble copper minerals should be put into solution and reprecipitated. It is sufficient if only a small percentage of the acid soluble copper minerals go into solution and are reprecipitated. lndeed since with appropriate conditioning as already illustrated a very substantial proportion of theseacid soluble minerals float in a completely alkaline circuit it is to be assumed that some dissolutionreprecipitation mechanism operates even at pHs well on the alkaline side of neutral. As will be apparent however from the examples already given and those to follow the floatability of the acid soluble copper minerals is maximized by the establishment of acid conditions in at least one stage of the total conditioning process. Further. the use of dispersing agents during such stage can be beneficial in ensuring that the effect of the treatment reaches the maximum proportion of particles in the pulp which are amenable to the treatment.

EXAMPLE XX Stage I A 1000 gram sample of the Lot 128 ore previously described was ground in a laboratory rod mill for 12 minutes at a pulp density of 50% solids with the addition of 2.0 lbs. per ton of H 80 Following grinding the pulp was transferred to a laboratory Fagergren where it was conditioned for minutes without any further reagent addition. The pH at the beginning and end of this conditioning period was 5.0.

Stage II 13.0 lbs per ton of Ca(OH) was added and conditioned for 5 minutes. The beginning pH was 12.20 and the final pH was 12.25. 7

Stage III 2.2 lbs per ton of H SO and 0.18 lbs per ton of Z6 were then added and the pulp was conditioned for 15 minutes. The beginning pH of this stage was 11.65 while that at the end was 1 1.40.

Stage IV 120 lbs per ton of sodium silicate were then added as a dispersing agent together with 4 drops of pine oil and the pulp was conditioned for a further 5 minutes. Throughout this stage the pH was 1 1.10.

The rougher concentrate was then floated during a period of 7 minutes with the addition of 0.02 lbs per ton of Z6 and 1 drop of pine oil. A small amount of sulphuric acid and Separan (a synthetic flocculant man- The rougher plus first cleaner tailings analysis was as follows:

'7! Weight 87.7 "/1 Total (u analysis of solids 0.040 '7! Acid Soluble Cu analysis of solids 0.019

'/( of total Cu in the solutions from the rougher plus first cleaner tailings 3.0 "/1 of original acid soluble copper in the solutions 12.5 '71 of total Cu in the first cleaner concentrate 93.3

EXAMPLES XXI, XXII and XXIII This series of tests used the same ore and circuitry as Example XX, to determine the effect of various alkali agents in combination with Ca( OH) and also omitting the acid addition in Stage III.

Stage II in all tests was taken as closely as possible to the same pH, that is, 11.65.

The following table shows the comparative results and the alkali agents used in Stage II.

Rougher 1st cleaner Alkali Agents Tailings Analysis lbs/ton Solids Solns. Example 70 7a of No. Ca() Na. ,CO;, NH OH NaOH Wt. Total Cu A.S.*Cu Total Cu XXI 8.2 4.0 88.0 0.063 0.028 2.0 XXII K6 2.0 87.6 0.055 0.018 4.3 XXIII 6.4 2.0 87.7 0.047 0.022 2.9

A5. Acid Soluble (u ufactured by Dow Chemical Co.) was added to the 45 rougher tailings and the latter was filtered.

The rougher concentrate was resuspended in a Den ver Cell and conditioned for 3 minutes with the addition of 0.02 lbs per ton of Z6 and 1.2 lbs per ton of sodium silicate. The thus conditioned pulp was then floated to cleanup.

The tailings were treated in the same manner as the rougher tailings and in each case there was less than 2 minutes total contact time in the cell and the filter together.

The head sample of this ore was 0.95% total Cu and 0.23% acid soluble Cu.

It will be noted that all combinations are effective with the sodium hydroxide being outstanding.

EXAMPLE XXIV In this example the same ore and circuitry was used as in Example XX. 3.2 lbs/ton of H was added to the grinding stage and 11.0 lbs/ton of Ca(OH) to Stage 11.

The rougher and first cleaner tailings were settled with a flocculant only, no sulphuric acid being used. The solutions at all stages were sampled and analyzed for copper. A

The following table shows the results obtained:

Stage pH Copper in Solution No. Description Begin End Mgs./I. Total Mgs.

. Begin I Grinding and conditioning 5.4 5.1 1.8 2.7

End

II Lime addition 12.2 12.2 Nil Nil Ill Collector addition 12.2 1 1.6 Nil Nil 1V Disperszmt addition 1 1.25 1 L2 Nil Nil End ol- Rougher Flotation Nil Nil End of First Cleaner Float Nil Nil The analysis of the rougher plus first cleaner tailings was 0.08% Total Cu, and 0.057% acid soluble Cu.-

The combined tailings weight was 90.5%

The recovery of the acid soluble copper in the first ples was employed using an all-alkaline circuit.

The solutions in each stage were sampled and analyzed for copper. Only traces of copper in one test showed up, and this was at a neutral pH.

The surprising phenomenon in the series of tests was the large reprecipitation of the copper dissolved in the tailings at a pH of 5.0 to 5.5 when re-circulated to the grinding circuit and transferred to the flotation cell.

The second surprising factor was the flotation of this copper following its re-precipitation with approxi- 5O sults over the prior described circuit, although from a plant operational point of view, it may, in some instances, be more desirable to add the acid to the grinding circuit only, or to the first conditioning step following the grinding circuit.

EXAMPLE XXVIII The same ore and circuitry as in the previous examcleaner f= was 5 The ore sample used contained 1.09% Total Cu and The surprisingly low amount of copper in solution in 0 3 Acid 1 1 Cu Stage I followed by the high percentage of acid soluble pp retalped m the first Gleaner concentrate were The typical metallurgical results were as follows: the outstanding features of this test, together with total Roushsr Plus Fir-$1 Cleaner Tuning-S Total Cu 0.] I re precipitation of the copper in Stage II that was in so- 10 9! Acid Suluhlc Cu M67 lution In Stage I. 7! Recovery Sulphide Cu in lst Clnr. Conct. 94.6 Recovery Acid Soluhlc Cu in lst Clnr. EXAMPLES XXV, XXVI and XXVII C n t. 83.0

In this series of tests, the same ore and circuitry was used as in Example XX. In the test prior to Example i recovefy of the acld Soluble copper m thls XXV 10 lbs on of H2804 was added to the rougher circuit is outstanding, particularly as the only collector tailingsand.2.8 lbs./ton of H 50 was added to the first used was the xifnthifte that "P f cleaner tailings. Following filtering of these tailings, m whlch i current Cnvem lOnal y 1000 of the Solution was recovered and added to cuit, has but little collecting effect on the family of acid the grinding stage of Example The Same soluble minerals. unless such modifiers as sulphidizing dure was followed with Examples XXVI and XXVII, agents are with the combined tailings solution from Example Two 9 the dlfferfmces my Over current XXV used in the grinding stage of Example XXVI, and conv'efimclnal pfacnce f much longer the combined tailings solution of Example XXVI used condition ng period of flotation, and a concentrated i the grinding Stage f Example XXV". use of dispersants in the latter stages of my pre- The objects of these tests was to determine whether conditioning Stagesor not h copper di l d i h ili ld By contrast in current conventional practice in the precipitate and float with the solids copper in th fl flotation of copper ores the pulp goes directly from the ti i i grinding circuit to the flotation cells (being adjusted to The solutions were sampled in all of the stages of the optimum pH for flotation either on the way there or three continuous tests. after arrival) and floated over a prolonged period of The following were results obtained: time with stage additions of the collector. In the pro- Solutions Solids Average pH Cu Mgs./l. Tailings Analysis Stage Example No. Example No. 71 T Cu "/1 AS. Cu No. xxv XXVI XXVII xxv XXVI xxvli xxv XXVI xxvii xxv XXVI XXVII I 5.3 5.3 5.2 8.9 7.8 5.7 11 11.x 11.8 1 1.8 0.0 0.0 1.2 111 1 1.7 1 1.65 1 L 0.0 00 0.0 IV 1 1.3 1 1.3 1 1.25 0.0 0.0 0.0 Rougher Plus lst Cleaner Tailings 144 0.060 0.057 0.051 0.026 0.021 0.017

cess of the invention even where the grinding stage is used as stage one it is necessary to use a minimum conditioning cycle following the grinding stage of 10 to 15 minutes for optimum results.

EXAMPLE XXIX Ore Lot 130 as previously described.

In this example sulphur dioxide has been used in Stage I to replace sulphuric acid, with approximately 60% of the S0 added to the grinding circuit and 40% to the conditioning cycle following grinding.

In Stage II the alkaline agent Ca(OH) and the collector Z6 were added simultaneously.

In Stage III the dispersing agent, Na SiO was added together with the frother, followed by the rougher float.

The first cleaner float was proceeded by a 3 minute conditioning cycle using 0.75 lbs. of NaQSiO per ton, and 0.025 lbs. of Z6 per ton.

The cleaner was insufficient to depress a satisfactory percentage of the remaining slirnes in the rougher concentrate. and a second stage of cleaning was required using a 2 minute pre-conditioning period with 0.5.

lbs/ton of Na SiO This illustration may be considered as a three stage pre-conditioning circuit to the rougher float. and an additional two stage dispersion circuit to the number 2 cleaner concentrate, to obtain satisfactory copper recovery in a concentrate containing an acceptably low clay content.

The following is the data on this example:

Stage I To grinding circuit 8.1 lbs/ton of S 12 minutes.

To conditioning 4.7 lbs./ton of S0 minutes.

pH end grinding 5.9

pH begin and end conditioning 4.8 and 5.4 respectively.

Stage II minutes conditioning with addition of 22 lbs/ton of Ca(OH) and 0.225 lbs. Z6/ton.

pl-I begin and end conditioning 11.7 and l 1.5 respectively.

Stage III 4 minutes conditioning with addition of 12.5 lbs/ton of Na SiO and two drops of pine oil Rougher Flotation.

The following results were obtained:

Rougher plus lst Cleaner Tailings 71 Weight 87.5 "/1 Total Cu analysis 0.068 "/r Acid Soluble Cu analysis 0.033 '7? Total Cu in lst cleaner concentrate 94.5 "/1 '7: Acid Soluble Cu in 1st cleaner concentrate 89.5 '71 4 Soluble Cu in solutions Nil. Rougher plus first plus second cleaner Tailings 7: Weight 92.5 '7: Total Cu in second cleaner concentrate 93.7

This test illustrated the following:

a. The satisfactory substitution of SO for H 80. in

my process.

b. The simultaneous addition of the alkaline agent and the collector following the acid conditioning stage.

c. The use of three stage circuit prior to rougher flotation and the ncessity of twostages of dispersion following the rougher flotation.

d. The complete re-precipitation of any acid soluble copper that had gone into solution in the acid cycle.

The high degree of flotability of the re-precipitated copper in conjunction with the sulphides.

EXAMPLE XXX Ore Lot 130 as previously described.

This example illustrates the use of a three stage conditioning circuit prior to rougher flotation, using S0 in place of H 50 The following is a description and data on the circuit used:

Grinding Circuit 12 minutes No reagents added pH at end 7.2

Stage I 10 minutes Conditioning with addition of 12.8

lbs/ton S0 and 0.625 lbs/ton of dispersant. lignin sulphonate.

pH begin 4.0, and end 4.1.

, Stage 11 As in Example XXIX Stage III As in Example XXIX Rougher Float Two cleaners as in Example XXIX.

At the end of Stage I the copper in solution was less than 5% of the total acid soluble copper and was totally precipitated in Stages I1 and III.

The metallurgy was practically identical to Example XXIX.

This test illustrated the following:

a. Using an acid circuit in the grinding stage would be eliminated, unless it was desirable from an operating point of view.

b. A dispersant could be used in the acid cycle where necessary without detrimental effects to the copper recovery.

EXAMPLE XXXI Ore Lot as previously described.

This example illustrates the use of a two stage conditioning system wherein a single acid stage is used and a single alkaline stage is used, prior to rougher flotation.

Two cleaners were used with dispersal of the pulp between each cleaner.

Grinding Circuit 12 minutes, no reagents added.

pH at end 7.2

Stage I 20 minutes conditioning, added 12.5 lbs/ton H SO and 0.625 lbs. per ton dispersant, tetra sodium pyrophosphate. ph begin 1.8, end 3.6.

Stage II 20 minute s conditioning, added 22.7 lbs/ton Ca- (OH) 0.25 lbs. /ton Na SiO and 2 drops of pine oil, at end prior to rougher flotation.

pH begin 11.9, end 11.4.

Rougher flotation followed by two stages of cleaning as in Examples XXIX and XXX.

The rougher plus first cleaner tailings was 86.4% by weight analyzing 0.057% total Cu and 0.023% acid sol uble Cu for a recovery in the first cleaner concentrate of 95.5% of the total copper and 94.5% of the acid soluble copper. There was no soluble copper in the tailings solutions.

This can be considered outstanding metallurgy for such a type of ore, particularly the flotation of the acid soluble copper.

The foregoing method of treating sulphuric acid soluble copper minerals whereby the copper in solution following an acid stage of the process is totally reprecipitated in a subsequent alkaline stage offers a very convenient and economic means of recovering sulphuric acid soluble copper contained in the tailings. It is possible of course to acidulate the tailings, putting the sulphuric acid soluble copper almost instantly into solution. filter, or thicken and employ the filtrate as makeup solution for an acid grinding stage at the head of my process. However, filtration is a relatively expensive operation and some mill operators prefer not to grind in decidedly acid circuts because of increased steel losses and possible scrap losses brought about by corrosion. Thus, an alternative is to pass the tailings to a tailings dump where they may be acidulated, and return to liquor overflow, which will contain dissolved copper, to an acid stage of conditioning which follows what may be a neutral or slightly alkaline grinding circuit. This alternative has outstanding merit in the case of operators having existing tailings dams and who already are recovering soluble copper from these dams by passing the tailings liquor through a special precipitation plant. By using the process of the invention the precipitation plant may be eliminated and the soluble copper from the tailings liquor or from any other source may be recovered directly by flotation.

What I claim as my invention:

1. A process for the flotation of copper values from copper ores containing copper sulphides and at least one of the sulphuric acid soluble copper minerals to separate said values from host rock materials containing clay or talcose materials, said process comprising: conditioning an appropriately prepared pulp of an ore containing such minerals in the presence of an alkaline agent and alkali metal xanthate at a pH of from about 9.5 to about 12 to produce heavy activation of said copper values; prior to flotation, subjecting the pulp to a final stage of agitation conditioning in the presence of a selected agent from the group consisting of dispersing agents and flocculating agents at a pH of from about 9.5 to about 12 to depress the host rock ingredients of the ore; and then subjecting the pulp containing both the copper sulphide values and the sulphuric acid soluble copper mineral values to froth flotation to produce a concentrate of said copper values.

2. A process according to claim 1 wherein the said flotation comprises a rougher float, and then a period of conditioning of the rougher concentrate by agitation with the addition of dispersing agent or flocculating agent to depress host rock ingredients, followed by a cleaner float.

3. A process according to claim 2 wherein the rougher concentrate is conditioned by agitation with lime at a pH in excess of 12 to about 12.35 prior to a cleaner float.

4. A process according to claim 1 in which the final stage of conditioning is continued for a period of at least minutes.

5. A process according to claim 1 wherein the ore contains molybdenum sulphide mineral and including at least one alkaline agitation conditioning step at a pH in the range of above 12.0 to about 12.3 prior to said final stage of conditioning, which final stage is conducted at a pH in the range of about 9.5 to about 12.0.

6. A process according to claim 1 wherein the alkali metal xanthate is potassium amyl xanthate.

7. A process according to claim 1 wherein a dispersing agent is used in said final stage of conditioning and the dispersing agent is sodium silicate.

8. A process according to claim 1 wherein the alkaline agent is lime, and the selected agent from the group consisting of dispersing agents and flocculating agents is a dispersing agent.

9. A process according to claim 1 wherein said appropriately prepared pulp of said ore is prepared by wet grinding said ore in the presence of an alkaline agent.

10. A process according to claim 1 wherein said conditioning an appropriately prepared pulp of an ore is substantially carried out by wet grinding said ore in the presence of an alkaline agent and an alkali metal xanthate at a pH of from about 9.5 to about 12.

11. A process according to claim 1 wherein said conditioning an appropriately prepared pulp of an ore is substantially carried out by wet grinding said ore in the presence of lime and alkali metal xanthate and the selected agent from the group consisting of dispersing agents and flocculating agents is a dispersing agent.

12. A process according to claim 1 wherein the alkaline agent is soda ash.

13. A process according to claim 1 wherein the alkaline agent is a combination of soda ash and lime and the final conditioning stage prior to flotation is conducted in the presence of a flocculating agent.

14. A process for the recovery by flotation of copper values from copper ores containing copper sulphides and at least one of the sulphuric acid soluble copper minerals said process comprising communiting said ore by wet grinding in the presence of calcium hydroxide and an alkali metal xanthate collecting agent to pro-- duce a prepared pulp of the ore, subjecting the said prepared pulp of the ore to at least one alkaline agitation conditioning stage within the pH range of about 9.5 to 12.0 in the presence of at least one alkaline agent selected from the group consisting of lime, calcium hydroxide, sodium carbonate, sodium hydroxide and ammonium hydroxide and in the presence of sufficient collector and in the presence of sufficient dispersing agent wherein the residence time of the pulp in at least one alkaline agitation conditioning stage is of sufficiently long duration to enable the effective recovery of copper values and effective depression of the waste non-sulphide host rock minerals in at least one subsequent flotation stage: and subsequently in the presence of a suitable frother subjecting the pulp to froth flotation to produce a flotation concentrate enriched in copper values and a tailing products low in copper values and high in waste non-sulphide host rock minerals.

15. A process as defined in claim 14 wherein a dispcrsant is present during wet grinding of the ore. 

1. A PROCESS FOR THE FLOTATION OF COPPER VALUES FROM COPPER ORES CONTAINING COPPER SULPHIDES AND AT LEAST ONE OF THE SULPHURIC ACID SOLUBLE COPPER MINERALS TO SEPARATE SAID VALUES FROM HOST ROCK MATERIALS CONTAINING CLAY OR TALCOSE MATERIALS, SAID PROCESS COMPRISING: CONDITIONING AN APPROPRIATELY PREPARED PULP OF AN ORE CONTAINING SUCH MINERALS IN THE PRESENCE OF AN ALKALINE AGENT AND ALKALI METAL XANTHATE AT A PH OF FROM ABOUT 9.5 TO ABOUT 12 TO PRODUCE HEAVY ACTIVATION OF SAID COPPER VALUES, PRIOR TO FLOTATION, SUBJECTING THE PULP TO A FINAL STAGE OF AGITATION CONDITIONING IN THE PRESENCE OF A SELECTED AGENT FROM THE GROUP CONSISTING OF DISPERSING AGENTS AND FLOCCULATING AGENTS AT A PH OF FROM ABOUT 9.5 TO ABOUT 12 TO DEPRESS THE HOST ROCK INGREDIENTS OF THEORE, AND THEN SUBJECTING THE PULP CONTAINING BOTH THE COPPER SULPHIDE VALUES AND THE SULPHURIC ACID SOLUBLE COPPER MINERAL VALUES TO FROTH FLOATION TO PRODUCE A CONCENTRATE OF SAID COPPER VALUES.
 2. A process according to claim 1 wherein the said flotation comprises a rougher float, and then a period of conditioning of the rougher concentrate by agitation with the addition of dispersing agent or flocculating agent to depress host rock ingredients, followed by a cleaner float.
 3. A process according to claim 2 wherein The rougher concentrate is conditioned by agitation with lime at a pH in excess of 12 to about 12.35 prior to a cleaner float.
 4. A process according to claim 1 in which the final stage of conditioning is continued for a period of at least 10 minutes.
 5. A process according to claim 1 wherein the ore contains molybdenum sulphide mineral and including at least one alkaline agitation conditioning step at a pH in the range of above 12.0 to about 12.3 prior to said final stage of conditioning, which final stage is conducted at a pH in the range of about 9.5 to about 12.0.
 6. A process according to claim 1 wherein the alkali metal xanthate is potassium amyl xanthate.
 7. A process according to claim 1 wherein a dispersing agent is used in said final stage of conditioning and the dispersing agent is sodium silicate.
 8. A process according to claim 1 wherein the alkaline agent is lime, and the selected agent from the group consisting of dispersing agents and flocculating agents is a dispersing agent.
 9. A process according to claim 1 wherein said appropriately prepared pulp of said ore is prepared by wet grinding said ore in the presence of an alkaline agent.
 10. A process according to claim 1 wherein said conditioning an appropriately prepared pulp of an ore is substantially carried out by wet grinding said ore in the presence of an alkaline agent and an alkali metal xanthate at a pH of from about 9.5 to about
 12. 11. A process according to claim 1 wherein said conditioning an appropriately prepared pulp of an ore is substantially carried out by wet grinding said ore in the presence of lime and alkali metal xanthate and the selected agent from the group consisting of dispersing agents and flocculating agents is a dispersing agent.
 12. A process according to claim 1 wherein the alkaline agent is soda ash.
 13. A process according to claim 1 wherein the alkaline agent is a combination of soda ash and lime and the final conditioning stage prior to flotation is conducted in the presence of a flocculating agent.
 14. A process for the recovery by flotation of copper values from copper ores containing copper sulphides and at least one of the sulphuric acid soluble copper minerals said process comprising communiting said ore by wet grinding in the presence of calcium hydroxide and an alkali metal xanthate collecting agent to produce a prepared pulp of the ore, subjecting the said prepared pulp of the ore to at least one alkaline agitation conditioning stage within the pH range of about 9.5 to 12.0 in the presence of at least one alkaline agent selected from the group consisting of lime, calcium hydroxide, sodium carbonate, sodium hydroxide and ammonium hydroxide and in the presence of sufficient collector and in the presence of sufficient dispersing agent wherein the residence time of the pulp in at least one alkaline agitation conditioning stage is of sufficiently long duration to enable the effective recovery of copper values and effective depression of the waste non-sulphide host rock minerals in at least one subsequent flotation stage: and subsequently in the presence of a suitable frother subjecting the pulp to froth flotation to produce a flotation concentrate enriched in copper values and a tailing products low in copper values and high in waste non-sulphide host rock minerals.
 15. A process as defined in claim 14 wherein a dispersant is present during wet grinding of the ore. 